Simulation of Gas Migration in Mines During Reversal Ventilation: A Case Study

Abstract

The objective of this study was to understand the characteristics of gas migration in a mine system network domain during a period of reversal ventilation. Combining field experiments with the TF1M3D simulation program, we analyzed gas migration and distribution during reversal ventilation in the JIU LI coal mine. The results showed that, after implementation of the airflow reversal process for the entire mine, the gas in the return roadways flowed back to the working face and accumulated with the gas emitted from the working face to form a gas concentration peak, after which the gas concentration gradually decreased in a stepwise manner and finally reached a stable state that was maintained until the end of the reversal ventilation. The peak gas concentration and the peak areas of the gas concentration curve during the airflow reversal were positively correlated with the time of airflow stoppage operation. The gas concentration peak affected the safety of the mine airflow reversal process; therefore, countermeasures and technical plans should be made in advance. The TF1M3D simulation results were consistent with the field experiment results.

1. Introduction

Underground mine fires threaten the safety of coal mines, and fires cause massive coal resource wastage and environmental pollution [1,2,3,4]. In the event of fires in mines, the priority is to take necessary and reasonable emergency relief measures to protect the lives of underground personnel. Fire and high-temperature smoke can disrupt the airflow of the ventilation system [5], and toxic gases from fires can spread and contaminate the area of production [6], which may lead to death by poisoning [7,8]. For example, on 27 September 2020, 16 people were killed and 42 injured in the “9–27” major fire accident at Songzao Coal Mine (Chongqing Energy Investment Yu New Energy Limited Company, China); 15 of the causalities were CO-poisoned. In the fire zone, high-temperature fires and smoke were present simultaneously, and the air volume decreased as the smoke accumulated. Such conditions are highly likely to lead to gas explosions in mines [9,10]. On 29 March 2013, the Babao coal company (Tonghua Mining Group of China Jilin coal group) saw a particularly serious gas explosion due to the spontaneous combustion of coal, causing the death of 36 people and injuring 12.

Regarding research on emergency rescue in the event of mine fires, numerous studies have been conducted on building mine ventilation safety information management systems to develop fire escape routes [11,12,13,14,15,16], establish airflow control systems to prevent fire deterioration [17,18,19], and construct life-saving facilities [20,21,22]. Ni et al. [15] established a comprehensive visual information management system for mine ventilation and safety based on a geographic information system. Wang et al. [16] used TEPS (Thunderhead Engineering PyroSim) to model a local mine ventilation system and develop fire smoke control measures. Charles [23] introduced a spatial navigation control system for a ventilation network, which was used to simulate the evacuation route during underground coal mine disasters. Wang et al. [17] used a mine ventilation variable-frequency system controller to regulate branch air volume and verified the practicality of variable-frequency regulation through experiments. Zhang et al. [6] determined the installation location of escape pods in the roadway based on numerical simulations and field experiment results. In addition, when fires occur in a mine, reversal ventilation of the mine can be achieved by reversing the operation of the ventilator [24], whereby the fire smoke airflow can be discharged from the intake shaft and limit the fire smoke airflow to the underground production areas. This is significant for reducing the scope of the disaster and effectively evacuating underground operators. The “Coal Mine Safety Regulations” of China [25] clearly stipulate that mines should conduct a reversal ventilation drill each year. The ability to perform reversal ventilation is an important technical indicator to determine whether a mine ventilation system can avoid potential disasters.

Mine reversal can be categorized as catastrophic ventilation. Under normal ventilation conditions, mine reversal only means the reversal of airflow in the ventilation network. In catastrophic ventilation, reversal not only involves the reversal of airflow, but must also reflect movement and change in atmospheric parameters such as temperature, humidity, and concentration of the gas components in the airflow of the mine ventilation network system, particularly movement and change in gas (CH₄) concentrations. The former is static and independent of time, whereas the latter is dynamic and temporal. The difference between the two is epitomized by normal ventilation and catastrophic ventilation, and mine reversal falls under catastrophic ventilation. The complexity of mine reversal varies, and in practice there are many elements, such as mine disaster reversals, that cannot be fully realized in reversal exercises, making it necessary to study them through computer simulations. For high-gas mines, mine reversal (drill) is a difficult task, and it is particularly significant to clarify changes in the gas concentration in the mine system during mine reversal. Therefore, reversal simulations study the unsteady motion of the transport–dispersion process of gas components in the ventilation network domain based on the ventilation network solution (airflow).

To understand the characteristics of gas migration in the network domain of a mine system during a period of reversal ventilation, this study combined reversal ventilation drill data with the TF1M3D simulation program of mine disaster ventilation developed by the authors to analyze the dynamic change characteristics of mine gas flow during reversal ventilation. The specific process is shown in Figure 1.

2. Materials and Methods

The coal mine is located in Jiaozuo City, Henan Province, China, and has a design production capacity of 0.9 Mt/a. The mine uses a vertical-shaft double-level up and down the mountain development mode, and the mine ventilation mode is two-wing diagonal ventilation. The mining method is inclined long wall mining. Figure 1 shows the ventilation system of the mine, which is fed by the main shaft and the secondary shaft and returned by the east shaft. The model of the east return air shaft fan is 2K60-4No.24, with a fan speeding of 740 r/min. The total air volume of the mine is 5391 m³/min, with a total resistance of 1888.4 Pa. The mine has two coal mining faces, namely 1101# located at the 11# coal seam and 15011# located at the 15# coal seam, and two tunneling faces, namely 15051# and 15061#; both the faces are located at the 15# coal seam. In Figure 2, the air volume (m³/s) of the roadway is indicated in blue, the arrows indicate the airflow direction, and the green numbers represent the label of the ventilation structure. According to the “Coal Mine Safety Regulations” [25]: (1) mines shall conduct a reversal ventilation drill each year, and (2) when there is a major change in the mine ventilation system, a reversal ventilation drill should be conducted. The reversal method for the JIU LI mine is to change the main fan from exhaust ventilation to forced ventilation.

The experiment was conducted during a reversal ventilation drill. According to coal mine safety regulations, the duration of reversal ventilation drill should not be less than 2 h, the airflow direction in the roadway must be changed within 10 min, and the air volume provided by the main fan during reversal ventilation should not be less than 40% of the normal air volume. Therefore, the total time of the drill was 120 min. During the drill, after 6 min of ventilation fan shutdown and reversal operation, the ventilation fan of the east shaft started to reverse the ventilation, thus achieving airflow reversal in the mine ventilation system. The inlet corner of the 1101# and 15011# working faces was selected as the observation site, and technicians who arrived at the observation site in advance used a JCB4 portable mine CH4 detector produced by Shandong Zhongzhong Intelligent Equipment Co., Ltd. (Shandong, China) to record the dynamic change process of the gas content during reversal ventilation. The gas measurement frequency was 60 s per time point. Three groups of data were measured at each time point, and the average value of the gas concentration was selected.

Figure 3 shows the dynamic changes in gas concentration recorded at two observation points. At the beginning of the reversal ventilation, as the ventilation fan shut down and reversal ventilation operated, the gas concentration in the original inlet roadway of the coal mining working face changed rapidly: the concentration at the inlet corner of the working face suddenly increased, and the first gas concentration peak appeared. After the first gas concentration peak at the working face, the gas concentration at the working face gradually trended toward a constant value that was maintained for a period of time. This was because the reversal operation changed the original ventilation direction of the mine system, resulting in the reverse flow of the gas accumulating in the original return air roadway into the working face, resulting in a slight increase in the gas concentration at the working face. At this point, the second gas peak occurred, which was smaller than the first one. During the period of mine reversal ventilation, a large amount of fresh air entered the mine system from the “East return air shaft” to dilute the residual gas in the mine return air roadway. Therefore, after the two gas peaks, the gas concentration at the inlet corner of the working face continued to decrease and reached a steady state again that was maintained until the end of the reversal ventilation drill.

3. Results

In the ventilation roadway, the equation of motion for the unsteady airflow can be expressed as:

$$P-\rho gS\sin \theta -{\displaystyle \frac{\lambda U}{2S^2}}\rho Q|Q|-{\zeta }_t{\displaystyle \frac{\rho Q^2}{2S}}=\rho S{\displaystyle \frac{dQ}{dt}}$$

(1)

$$r_t={\xi }_t\cdot {\displaystyle \frac{\lambda }{8}}\cdot {\displaystyle \frac{U}{S^3}}$$

(2)

where P is the average wind pressure at the roadway section, expressed in Pa units; ρ is the average airflow density at the roadway section, expressed in kg/m³; Q is the air volume, expressed in m³/s; g is the gravitational acceleration, expressed inf m/s²; θ is the inclination of the roadway; U is the perimeter of the roadway, expressed in m; S is the sectional area of the roadway, expressed in m²; τ is the time variable, expressed in s; r_t is the geometric windage per unit length of the roadway, expressed in 1/m⁵; λ is the resistance coefficient of the Darcy–Weisbach formula, and is dimensionless; And ζ_t is the proportional coefficient of the local resistance of ventilation, and is dimensionless. Generally, ζ_t ranges from 1.05 to 1.15.

The convection–dispersion equation of the gas in the airflow of a single ventilation roadway during the period of reversal ventilation can be expressed as:

$$\begin{gathered} {\displaystyle \frac{\partial C}{\partial t}}+v\cdot {\displaystyle \frac{\partial C}{\partial x}}=E_x\cdot {\displaystyle \frac{{\partial }^{2}C}{\partial x^2}}+W_s \\ \mathrm{Boundary}\mathrm{condition}:C_i{\mid }_{i=\mathsf{\Gamma }}=0;C{\mid }_{x=0}=C_i \end{gathered}$$

(3)

where C is the volume percentage of the average gas concentration in the roadway section, expressed as %; v is the average wind speed of the roadway, expressed in m/s; E_x is the longitudinal mechanical dispersion of the roadway airflow, expressed in m²/s; α is the roadway coefficient of the frictional resistance, expressed in N·s²/m⁴; W_S is the gas effusion (source) intensity at the working face, expressed in m³/s; Γ is the mine air inlet wellhead boundary; i is the starting point of a branch of the mine; and C_i is the gas concentration at the nodal point.

The gas effusion is a typical external ventilation source; therefore, it was treated as a wind source in the active ventilation network. Accordingly, the mass balance equation for the wind network in a certain time step is as follows:

$$A\cdot M=A^{*}\cdot W_S+D$$

(4)

where A = [a_ij]_(m-1)×n is the basic node incidence matrix. When the i node is the starting point of the j branch, a_ij = 1; when the i node is the end point of the j branch, a_ij = −1; otherwise, a_ij = 0. M = [M_j]_n×1 is the ventilation mass flow vector, M_j is the mass wind flow of the branch j, kg/s, where M_j = ρ_j Q_j; Q_j is the volumetric airflow (air volume) of the branch j, m³/s; ρj is the density of the branch wind flow of j, kg/m³; W_S = [W_j]_n×1 is the branch source term (vector), where W_S represents the branch weak source, W_j is the branch wind source, the mass flow of the source term (gas effusion or gas produced by fire) on branch j, kg/s; D = [D_i]_(m-1)×1 is the node source term (vector), D_i is the node wind source, the mass flow rate of the source term (gas effusion or gas produced by fire) at node i, kg/s. Here, A* = [a*_ij]_(m-1)×n is the basic node aggregation matrix. When a_ij = −1, a*_ij = 1; otherwise, a*_ij = 0.

The major difference between an active ventilation network and an ordinary ventilation network is that the right-hand side of the mass balance equation in the active ventilation network is not zero, but instead includes the source and sink terms. For branch weak sources, Equation (4) can be rewritten as:

$$\begin{gathered} {\displaystyle \sum _{j=1}^n{a}_{ij}\cdot M_j=}{\displaystyle \sum _{j=1}^{n}a{*}_{ij}\cdot W_j}+D_i \\ (i=1,2\dots n) \end{gathered}$$

(5)

When the branch weak source is ignored, or when the branch strong source can be directly converted to a node wind source, we have the following equation:

$${\displaystyle \sum _{j=1}^n{a}_{ij}\cdot M_j=}{D}_i$$

(6)

From Equation (4), the expression is complete in the form of the equation. Therefore, in a broad sense, the active ventilation network is the general form of the network, and the ordinary network (homogeneous equation) is only a special case of the general form.

The energy (the wind pressure balance equation) balance equation for an active ventilation network is as follows:

$$B\cdot H=B\cdot H_f+B\cdot P_e+B\cdot \delta$$

(7)

where B is the basic loop matrix; H is the wind pressure vector, Pa; and P_e = [Pe,j]_n×1 is the position pressure difference vector, where P_e is the position pressure difference of the j branch, Pa. At the same time, P_e,j = (Z_j,2–Z_j,1) ρ_j g, where Z_j,1 and Z_j,2 are the elevations of the start and end nodes of branch j, respectively, expressed as m. δ = [δ_j]_n×1 is the airflow unsteady term vector.

$${\delta }_j={\rho }_j\cdot {\displaystyle \frac{L_i}{S_i\cdot \mathsf{\Delta }\tau }}\cdot \left(Q_j^{\left(0\right)}-Q_j\right)$$

(8)

where L_j is the length of j branch with the unit of m, Δτ is the step of time with the unit of s, and Q_j⁽⁰⁾ is the initial value of the air volume (air volume value of the previous time) of j branch with the unit of m³/s.

The reversal ventilation start process is not a simple airflow reversal, but generally goes through ventilator shutdown, reversal ventilation operation, power transfer and airflow start, and unsteady change. During this short duration, the airflow in the mine is in a short near-stagnant state, which can lead to gas accumulation. Therefore, the “Coal Mine Safety Regulations” stipulate that this duration should not exceed 10 min. In this study, the reversal ventilation start process was divided into three steps: (1) ventilator shutdown, (2) reversal ventilation operation (stagnation for a period of time), and (3) completion of mine airflow reversal. The objective of the simulation was to describe the process of airflow change and subsequent gas accumulation change in the mine ventilation system during this process. The simulation calculation of Equations (1) to (8) was based on the mine ventilation network solution to complete computer simulation using numerical calculation methods. TF1M(3D) is a 3D mine disaster ventilation simulation program developed by the authors based on the source ventilation network theory [26]; it is used to solve the mine ventilation network problems and to simulate transport of the gas component of the airflow, temperature, and environmental parameters during mine disaster calculation. The program allows analysis and optimization of an actual mine ventilation system.

Taking JIU LI mine as the engineering background, the mine was simulated using the TF1M3D simulation program for reversal ventilation. The constructed 3D model structure data and ventilation system basic data and calculation parameters were first inputted to TF1M3D to generate the mine ventilation system model. The ventilation simulation calculation included 193 coordinate points, 242 roadway sides, and 21 additional resistance doors. Figure 4 shows the gas distribution in the mine before reversal ventilation, where the color scale on the right side of the figure represents the gas concentration, and the red numbers represent the gas effusion rate, expressed in m³/min. In the figure, the red numbers represent gas emissions at each measuring point, green numbers represent the air volume of the roadway (m³/s), and the arrows indicate the airflow direction.

During the period of mine reversal ventilation, the ventilator was switched from exhaust ventilation to forced ventilation, and the speed of ventilator was adjusted from n = 740 r/min to n = 340 r/min in the simulation to reflect the reduced ventilation capacity. The total time of the reversal ventilation in the simulation was set to 2 h.

Table 1. Ventilation and gas information before and after reversal ventilation.

Working Face	Normal Ventilation Period			Reversal Ventilation Period
	Air Volume (m3·s−1)	Gas Emission Rate (m3·min−1)	Gas Concentration (%)	Air Volume (m3·s−1)	Gas Emission Rate (m3·min−1)	Gas Concentration (%)
1101#	12.41	5.51	0.73	3.486	2.76	0.90
15011#	12.86	5.32	0.69	3.402	2.66	0.85

presents the ventilation and gas information before and after reversal ventilation.

The TF1M3D simulation program provided 300 dynamic images of the gas distribution in the entire mine system after reversal ventilation. Four of them are shown here to represent the dynamic changes in gas distribution in the mine after reversal ventilation. Figure 5 shows the dynamic changes in gas distribution in the mine after reversal ventilation. The changes in gas concentration at the working face during the period of reversal ventilation were as follows: (1) when reversal ventilation started, the gas concentration increased significantly, and (2) as reversal ventilation of the mine proceeded, the gas concentration gradually decreased and finally stabilized at a certain value.

Comparison of Figure 3 and Figure 4 shows that, during the first 300 s of the shutdown and reversal operation, natural ventilation pressure played a major role [27]. Airflow in the mine was largely in a sluggish state at this time, and the gas at the return shaft appears to have returned, while gas effusion from the working face accumulated at the working face. At 300 s, the mine completed reversal of the ventilation system airflow, and at the same time, the reverse flow of gas in the original return air ducts of the 1101# and 15011# working faces returned to the coal mining face, and the gas concentration at the corner of the return air increased. The reverse flow of the gas and the gas effusion from the working face combined at the working face and then flowed out from the inlet of the working face. At 852 s, the c gas concentration at the corner of the inlet of the 15011# coal mining face was 1.6%. At 948 s, the gas concentration at the 1101# coal mining face was 1.64%. Due to the lower efficiency of the ventilation fan (lower air volume) after reversal ventilation, the gas concentration in the mine ventilation system was higher than that during the normal ventilation period until 7200 s, the end of the reversal ventilation drill.

The air volume of the mine system decreased during the period of reverse air stoppage; therefore, there was a certain accumulation of gas. The longer the airflow was stopped, the more the gas accumulated, with a higher peak gas concentration during the period of ventilation stoppage. To prevent the gas concentration from increasing or even exceeding the limit, reversal ventilation is generally carried out when the working face is shut down or maintained, so as to keep the absolute gas outflow at the lowest possible level. When production is shut down or maintained, there is no gas outflow from the mined coal wall; therefore, the gas outflow from the coal mining face induced by reversal ventilation should be reduced.

Figure 6 shows the gas concentration change curve at the corner of the inlet of each coal mining working face, as generated by TF1M3D. As shown in Figure 6, the gas concentration in the original inlet of the coal mining working face changed quickly at the beginning of the reversal ventilation, and the gas concentration in the inlet of the working face increased suddenly, which was the result of the gas outflow at the working face. The peak was the result of gas outflow in the superimposed section of the working face. After the peak of the superimposed gas at the working face, there was a period of stability, which was influenced by the gas concentration in the total return airflow roadway and led to a peak in the gas sub-stack, after which it reached a stable state again that lasted until the end of the reversal ventilation. Comparing the measured results with the simulation results in Figure 6, the measured peak gas concentration at the 1101# coal face was 1.62%, the gas concentration in the steady state was 0.94%, and the simulated values were 1.64% and 0.90%, respectively. The errors were 1.23% and 2.13%; the measured peak gas concentration at the 1501# coal face was 1.60%, the gas concentration in the steady state was 0.85%, and the simulated values were 1.60% and 0.95%, respectively, which were the same as the measured values. The error between the simulated and measured results was less than 3%, and the simulation results obtained using TF1M3D were consistent with the measured results.

A short period of airflow stoppage after the start of reversal ventilation can lead to gas accumulation, which in turn affects the magnitude of the gas concentration peak. Figure 7 shows the relationship between stoppage time and gas concentration. Comparing the simulation results at 10 min, 6 min, and with (ideally) nonstop airflow stoppage, we found that the peak gas concentrations were 1.93%, 1.64%, and 1.48%, respectively, at the corner of the inlet of the 1101# working face. The peak values of the gas concentration with operation times of 10 min and 6 min were 1.30 times and 1.11 times that of nonstop airflow (under ideal conditions), respectively; the peak gas concentrations were 1.89%, 1.60%, and 1.40%, respectively, at the corner of the inlet of the 15011# working face. The peak value of the gas concentration with operation times of 10 min and 6 min were 1.35 times and 1.14 times that of nonstop airflow (under ideal conditions), respectively. Figure 8 shows the areas of each gas concentration peak in Figure 7. The longer the airflow stoppage operation time, the larger the area of the gas concentration peak, indicating that the length of airflow stoppage affected gas accumulation during reversal ventilation. The increase in the area of the gas concentration peak for the 1101# coal face at 10 min and 6 min was 1.22 times and 1.26 times, respectively, and that for the 15011# coal face was 1.35 times and 1.51 times, respectively. Because the airflow volume in the mine system decreased during reversal ventilation, there was some gas accumulation. Therefore, in addition to implementing the Coal Mine Safety Regulations, which stipulate that reversal ventilation must be completed within 10 min, it is necessary to minimize reversal ventilation downtime during the period of reversal ventilation and ensure that the ventilator reversal device is intact and effective during normal ventilation.

The error between the simulation results obtained by TF1M3D and the measured results was less than 3%. The simulation results were consistent with the typical characteristics of airflow reversal at the actual site. These results can help establish an analysis platform for reversal ventilation of mines and a theoretical prognosis for reversal ventilation operation.

4. Conclusions

(1) During reversal ventilation of mines, the gas concentration at the working face inlet roadway (the same as the return roadway during reversal ventilation) exhibits a general increase accompanied by two peaks. The reasons are as follows: (1) brief gas accumulation caused by stagnation of the airflow during reversal ventilation; (2) unfavorable gas discharge caused by a decrease in the air volume after reversal ventilation; and (3) reversal ventilation of the airflow containing gas from the original return airway at the working face, which converges with gas effusion from the working face.

(2) During the reverse airflow period, short-term stagnation of airflow after reverse airflow is initiated can lead to the accumulation of gas. The stagnation time is positively correlated with the peak concentration of gas. When the stagnation time extends from 0 min to 10 min, the peak concentration of gas increases by approximately 1.30 to 1.35 times, which is highly likely to increase gas concentrations past safe limits and cause significant safety hazards. Therefore, in addition to completing reverse airflow within 10 min as required by the “Coal Mine Safety Regulations”, the reverse airflow shutdown time should be shortened as much as possible. It is necessary to ensure that the reverse airflow fan is in good condition and effective on a daily basis.

(3) During the period of reversal ventilation, to avoid gas overrun, response measures and technical plans for adjusting production should be implemented. Because the gas concentration at the inlet of the working face was lower than that at the return during the period of normal ventilation, the gas concentration at the return should be detected as a matter of priority and continuously monitored during airflow reversal to avoid gas overrun during reversal ventilation.

(4) The error between the simulation results obtained by TF1M3D and the measured results was less than 3%. The simulation results were consistent with the typical characteristics of the airflow reversal at the actual site, which can provide a theoretical prediction for reversal ventilation in mines.